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IC 9014 



Bureau of Mines Information Circular/1985 



Remedial and Strata Replacement 
Techniques on Longwall Faces 

A Report on the State of the Art 
By Robert S. Dalzell and Ernest A. Curth 




UNITED STATES DEPARTMENT OF THE INTERIOR 



75i 

"AMINES 75TH A^ 



Information Circular 9014 

1 1 



Remedial and Strata Replacement 
Techniques on Longwall Faces 

A Report on the State of the Art 



By Robert S. Dalzell and Ernest A. Curth 




UNITED STATES DEPARTMENT OF THE INTERIOR 
Donald Paul Hodel, Secretary 

BUREAU OF MINES 
Robert C. Horton, Director 







Library of Congress Cataloging in Publication Data: 



Dalzell, Robert S 

Remedial and strata replacement techniques on longwall faces. 

(Information circular ; 9014) 

Bibliography: p. 27-29. 

Supt. of Docs, no.: I 28.27:9014. 

1. Ground control (Mining). 2. Coal mines and mining— Safety mea- 
sures. I. Curth, Ernest A. II. Title. III. Series: Information circular 
(United States. Bureau of Mines) ; 9014. 



TN295.U4 [TN2881 622s [622'. 334] 84-600294 



CONTENTS 

Page 

Abstract 1 

Introduction 2 

Acknowledgments 4 

Wire meshing 4 

Chemical rock stabilization 10 

Conventional cavity filling 15 

Strain-absorbent modules 15 

Inf latables 17 

Chemical foam 19 

Monolithic fills 22 

Planned Bureau research 26 

Summary and conclusions 26 

References 27 

ILLUSTRATIONS 

1. Shield face with chainless haulage rack 2 

2 . Thin-seam shield face with plow 3 

3. A longwall face in a void 3 

4 . Mechanical wire-mesh applicator 5 

5 . Wire-mesh screen on canopies and at rear of chocks 7 

6 . Wire mesh laid on canopies 8 

7 . Shield recovery under wire mesh 9 

8. Recovery area secured with wire mesh, bridge boards, and roof bolts 9 

9 . Recovery area secured by cribs under wire mesh 10 

10. Coal face stabilization using resin cartridges 12 

11. Roof stabilization by polyurethane injection 12 

12. Area stabilized by polyurethane injection during shield recovery 14 

13. Strain absorbent modules 16 

14. SAM tested under 30-kN load 17 

1 5 . Inf latables 18 

16. Dunnage bag inflated to 12 kPa supporting a 1,270-kg vehicle 19 

17. Isoschaum foam application 20 

18. Foam-filled vent tubes on base of forepoles and split bars 20 

19. Cavity filling with foam 21 

20. Cavity over coal face hading forward 23 

21. Aqua-packing system 24 

22. Strata replacement in a high and wide cavity 24 

23. High cavity on shield face 25 

24. Strata replacement scheme used to fill cavity shown in figure 23 25 

TABLES 

1. Incidence of longwall ground-control-related accidents, 1977-82 2 

2. Reactive characteristics of polyurethane systems 11 

3. Comparison of various materials for filling a 10-m 3 cavity 21 



L, 



i 





UNIT OF MEASURE 


ABBREVIATIONS 


USED 


IN THIS REPORT 


cm 


centimeter 






m 2 


square meter 


ft 


foot 






m 3 


cubic meter 


h 


hour 






min 


minute 


in 


inch 






mm 


millimeter 


kg 


kilogram 






pet 


percent 


kg/L 


kilogram per 


liter 


ppm 


part per million 


kN 


kilonewton 






s 


second 


kPa 


kilopascal 






t 


metric ton 


L 


liter 






t/h 


metric ton per hour 


lb 


pound 






yr 


year 


m 


meter 











REMEDIAL AND STRATA REPLACEMENT TECHNIQUES ON LONGWALL FACES 

A Report on the State of the Art 

By Robert S. Dalzell ' and Ernest A. Curth 2 



ABSTRACT 

Following the introduction of roof shields to the U.S. longwall min- 
ing scene in 1975, the number of shield faces (faces supported by 
shields) has increased steadily, and the occurrence of accidents asso- 
ciated with ground control inadequacies at longwall operations has 
shown a downward trend. However, ground control problems in longwall 
mining still exist. New mining systems, however efficient, have run 
afoul of roof control. Shield faces have suffered extended stoppages 
because of roof cavities. 

In this report, the Bureau of Mines addresses the problem of longwall 
roof control. The report presents a review of foreign remedial and 
strata replacement techniques for longwall faces and evaluates their 
potential for domestic application. The technology reviewed includes 
meshing, chemical rock stabilization, cribbing, strain-absorbent mod- 
ules, inflatables, synthetic foams, and monolithic fills. 

1 Mining engineer, Pittsburgh Research Center, Bureau of Mines, Pittsburgh, PA (now 
with Jim Walter Resources, Inc., Brookwood, AL) . 

2 Mining engineer, Pittsburgh Research Center, Bureau of Mines, Pittsburgh, PA. 



INTRODUCTION 



Table 1 shows the recent decrease in 
the number of accidents associated with 
longwall ground control. The advent of 
shields (figs. 1-2) to the U.S. longwall 
mining scene in the seventies and their 
predominance in the eighties probably ac- 
counts for much of the accident reduc- 
tion. The development of roof shields in 
West Germany preceded and paralleled the 
rapid adoption of roof shields by U.S. 
miners. In 1982, 85 pet of the produc- 
tion from West German coal mines came 
from shield faces ( 1_) . 3 

Shields have several safety and produc- 
tivity advantages over chocks. Shields 
provide — 

• A sheltered working space requiring 
minimum cleanup work. 

■^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this report. 



TABLE 1. - Incidence of longwall ground- 
control-related accidents, 1977-82 











Ground- 


Year' 


Total 
faces 


Shield i 


faces 


control- 




Number 


pet 


related 










accidents 2 


1977 (2) 


77 


15 


19 


143 


1978 


NC 


NC 


NC 


59 


1979 (3) 


91 


40 


43 


60 


1980 (4) 


105 


57 


54 


89 


1981.... 


NC 


NC 


NC 


73 


1982 (5) 


112 


93 


83 


87 


1983 (6) 


118 


99 


84 


71 



NC No census available. 

'Underlined numbers in parentheses re- 
fer to censuses of longwall faces listed 
in the references at the end of the 
report. 

2 Source: Health and Safety Analysis 
Center, MSHA, U.S. Department of Labor. 




FIGURE 1. - Shield face with chainless haulage rack. 




FIGURE 2. - Thin-seam shield face with plow. 



• Structural stability that allows ad- 
vancing without delay, even with 
brushing roof contact. 

• Lemniscate gear that provides a nar- 
row and nearly equal span of exposed 
roof ahead of the canopy over the 
entire vertical range. 



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A less severe 
(West German mi 
sion to describ 
above powered 
by frequent se 
Shields take 
from gob to 
canopies are 
canopies . 



"trampel" effect, 
ners use this expres- 
e roof deterioration 
roof supports caused 
tting and lowering.) 
fewer steps walking 
ace, because shield 
shorter than chock 



However, a persisting problem is roof 
cavities above the roof supports. These 
cavities increase both in height and ex- 
panse, and the problem is exacerbated 
because the wide canopies of shields re- 
strict access to the cavities for reme- 
dial work. Roof rock deterioration, if 



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FIGURE 3. - A longwall face in a void. 

untreated, can lead to rock breaking for- 
ward in front of the canopies (fig. 3). 
Such deterioration must be prevented. 
Otherwise, long delays to restore roof 
control will impair productivity and in- 
duce safety hazards. 



Results of studies on 132 longwall 
faces indicated the severity of the 
strata deterioration problem in West 
German mines. Cavities more than 50 cm 
high over more than 10 pet of the face 
length were observed on numerous long- 
walls and were found to be related to 
the support capacity and the span of 
exposed roof ahead of the canopies 
(7). 

Great Britain's National Coal Board 
(NCB) was confronted with the same 
problem, uncontrolled cavities above 
and ahead of the shields , when it in- 
troduced shields to U.K. mines in re- 
cent years. Task forces appointed to 
research remedial techniques conducted 
operations studies on numerous faces 
(8). 



The Bureau of Mines recognizes the 
need for remedial and strata replacement 
techniques for longwall faces in the 
United States and is reviewing foreign 
technology in order to evaluate its po- 
tential for domestic application. The 
parameters being considered are minimum 
exposure of workers to hazardous roof, 
effectiveness, speed, simplicity, and 
cost. The goals of the Bureau's research 
on remedial and strata replacement tech- 
niques are long range: to enhance safe- 
ty, productivity, and resource recovery 
where extraction is vulnerable to ground 
hazards. Future research efforts will 
be directed to fitting remedial systems 
into the mining cycle, the development 
of chemical foam without harmful side 
effects to miners, and the technology of 
monolithic fills. 



ACKNOWLEDGMENTS 



The authors are indebted to Kirk 
McCabe, product manager, Mobay Chemical 
Corp., Pittsburgh, PA, and Derek Tuke, 
president, Micon Services, Inc., Pitts- 
burgh, PA, for valuable information on 
chemical strata consolidation. The au- 
thors acknowledge permission granted by 



Gluckauf to use figures 3, 10-11, and 22; 
by Tinsley Wire (Sheffield) Ltd. to use 
figures 4-5; by the NCB, including the 
Mining Research and Development Estab- 
lishment, to use figures 13-15, and 17- 
19; and by The Mining Engineer to use 
figures 20-21, and 23-24. 



WIRE MESHING 



In the early days of powered roof sup- 
port, caved rock flushing in from the gob 
and debris of friable stratum dropping 
out between the canopies created a hazard 
in the working area and required labor- 
intensive cleaning of the operating space 
so the roof supports could be advanced. 
The "trampel" effect (roof deterioration 
above powered roof support) was, and con- 
tinues to be, particularly harmful. 

Prior to the advent of roof shields, 
European miners stretched wire mesh of 
high strength and adequate elasticity 
over the canopies to hold the fragile 
roof material and thus prevent the 
harmful ingress of debris into the travel 
space. A report on West German mining 
technology in 1971 described four faces 
where the roof space was covered with 
wire mesh. On one face, the wire mesh 
was paper backed to shelter the face area 
from dust originating from the gob during 
advance of the roof supports (9). In 



1973, West German statistics on longwall 
operations indicated that 31 pet of a 
total of 159 active faces with powered 
supports required the use of wire mesh 
(10). 

Two types of wire mesh were in use, or- 
dinary wire netting of 50-mm mesh and 
1.5- to 2.5-mm wire (16 to 12 SWG), which 
is readily available, and welded wire 
mesh with 40-mm mesh and 1.5-mm wire (16 
SWG). Since it is welded at the knots, 
welded mesh is rather stiff, and there- 
fore comes in sections only 1.5m wide 
and 25 m long. Welded mesh weighs less 
than wire netting of comparable strength 
and suffers less sag because of its 
stiffness. 

Threading wire mesh over the canopies 
and splicing the sections together is 
highly labor intensive. Special steel 
clips are used, and West German mining 
research has developed a plier-like tool, 



combined with a clip dispenser, to con- 
nect the sections together. To reduce 
the time required for splicing, 4-m-wide 
sections instead of the usual 1.5-m ones 
were fabricated and folded into 20-cm- 
diam rolls to facilitate transportation, 
storage, and handling on the face. 

In Great Britain, the Mining Research 
and Development Establishment developed 
mechanical mesh laying to reduce the 
time-consuming, arduous, and hazardous 
manual effort; and a report from 1980 
indicates that this technique was used in 
17 collieries on 25 faces (11). Face 
manpower was reduced and performance 
greatly improved. The usual practice of 
leaving a substantial layer of roof coal 



to hold a fragile rock stratum could be 
abandoned, and instead the shearer cut up 
to the immediate roof, resulting in bet- 
ter resource recovery. 

A 50- by 50-mm welded wire mesh with 
1.5-mm wire (16 SWG) is recommended for 
mechanical mesh laying. A section (75 m 
long and 70 cm wide) is supplied in a 48- 
cm-diam roll that weighs 32 kg. 

The latest version of a mechanical wire 
mesh applicator was developed at the 
Blidworth Mine in northern Nottingham- 
shire County, U.K. (12). It consists of 
a holder framework that accepts a roll of 
wire mesh (fig. 4). The framework, which 
is attached to a shearer cowl, presses 




FIGURE 4. - Mechanical wire-mesh applicator. 



the wire-mesh roll to the roof by spring- 
loaded arms so the roll can follow the 
outline of the roof. The spring-loaded 
arms also allow enough yield to keep 
damage from falling rock to a minimum. 
One end of the wire mesh is secured to 
the face-end roof support, and, as the 
shearer traverses the face, the wire mesh 
is pulled off the roll with very little 
sag. The roof supports can then be ad- 
vanced under the mesh, and the shearer 
can cut without interference. 

To replace a wire-mesh roll, the 
shearer operator raises the ranging drum 
that lets the cowl turn and drop down- 
ward; the face crew can then reload the 
mesh applicator. To resume mesh appli- 
cation, the operator maneuvers the cowl 
back into the operating position, press- 
ing the wire-mesh roll against the roof. 
The face workers then clip the wire 
mesh from the new roll to the already- 
installed mesh. The mesh applicator can 
be easily detached from and remounted to 
the cowl at the face ends , when the cowls 
are reversed. 

Handling the wire mesh and splicing the 
sections together still requires a great 
deal of manual effort. But the hazardous 
part of the work, namely, threading the 
wire mesh over the canopies with the 
miners exposed to unsupported roof, is 
eliminated by mechanical mesh laying. 
Also, the expense for manpower to remove 
debris from the supported area is saved. 

Of course, roof shields are very effec- 
tive in sheltering the working space from 
debris and dust. Their introduction to 
NCB mines in the late seventies elimi- 
nated the need for mesh laying on faces 
so supported except during salvage opera- 
tions (shield recovery). Shield recovery 
poses an unusual problem, because part of 
the structure is under caved roof. Wire 
mesh, applied prior to face termination, 
forms a protective blanket under which 
recovery can take place unimpeded by de- 
bris from the caved roof rock. 

In the United States, mesh laying was 
practiced on a longwall face in the York 
Canyon Coalbed in New Mexico in 1974. 



The coalbed was 2 m thick. A wide fault 
dissected the face and had to be blasted. 
The roof was stabilized with 1.8-m resin 
bolts, and wire mesh was stretched over 
the canopies of the four-leg 4,450-kN 
chocks to keep debris from dropping be- 
tween the roof supports (fig. 5). Pro- 
ductivity under these conditions was very 
low and did not improve as long as the 
fault intersected the face. 

At present, wire mesh is applied to 
U.S. longwall faces only during the sal- 
vage phase of operations that use 
shields. A protective mat is formed by 
threading imported wire mesh or domestic 
cyclone fence over the canopies, begin- 
ning 10 shears or 7.5 m from the terminal 
position of the face. Imported wire mesh 
is a semitwist 40- by 40-mm square mesh 
of 1.5-mm wire that is very flexible and 
easily handled in the confined longwall 
space; however, it is expensive and not 
always available. The domestic cyclone- 
fence-type netting is a 50- by 50-mm 
square mesh of 3- or 3.8-mm wire (13) . 

The wire mesh is laid by unrolling the 
sections either parallel to the face or 
in the direction of face advance (14) 
(figs. 6-7). The sections are spliced 
together with steel clips, leaving 15-cm 
overlaps. The wire-mesh mat is rein- 
forced with 5- or 15-mm wire rope, which 
is stretched parallel to the face from 
gateroad to gateroad and anchored to 
roof bolts. One wire rope is installed 
for each shear at 0.75-m intervals. The 
recovery space is also secured with 
bridge boards and roof bolts (fig. 8). 
In thick seams , meshing the face prevents 
sloughing of the coal (15) . 

During the recovery phase, the salvage 
crew builds cribs in place of each re- 
moved shield while the wire mesh is held 
up by hydraulic props (fig. 9). 

The following example illustrates how 
the quantity of wire mesh needed may 
be calculated: A blanket of wire mesh 
is to be installed on a 150-m face prior 
to shield recovery, taking in 10 strips 
(shears) or 7.5 m. Using 1.5-m-wide 
rolls of wire mesh, and leaving 15-cm 




FIGURE 5. • Wire-mesh screen on canopies and at rear of chocks. 










i 










FIGURE 6. - Wire mesh laid on canopies. 




FIGURE 7. - Shield recovery under wire mesh. 



5-ft resin 
bolts- 




FIGURE 8. - Recovery area secured with wire mesh, bridge boards, and roof bolts. 



10 



** 




FIGURE 9. - Recovery area secured by cribs 
under wire mesh. 

overlaps, six lengths of mesh 150 m long 
are needed to cover the roof area. The 
total quantity of wire mesh is 



Compared with manual mesh application, 
mechanized mesh laying, as practiced in 
the United Kingdom, may reduce hazards 
and costs in the shield recovery process. 
A longwall face, supported by shields and 
mined by a double-ended ranging-drum 
shearer, has a potential for an average 
daily output of 3,120 t of coal (from 
three production shifts). Each 0.75-m 
strip on a 150-m face extracting 1.80 m 
of coal contains 260 t of coal, so that 
four strips are needed to fulfill the set 
shift production goal. 

Ten strips without mesh blanketing 
could be mined in 2.5 shifts (10 t 4). 
Manual mesh laying would reduce shift 
performance to an estimated 1.5 strips, 
meaning that 10 strips could be mined in 
6.5 shifts. Thus, four additional shifts 
would be needed to accomplish the mesh 
laying, resulting in a revenue shortfall 
of $152,000 (4,160 t x $36.50), given the 
average coal output for four production 
shifts and the current cost of utility 
fuel. The opportunity loss for 1 day of 
stoppage is $7,000, with $8 million in- 
vested and a 20-pct return on investment, 
in addition to $6,000 for face labor. 
The cost for 1-1/3 days of work interrup- 
tion (four shifts) would be as follows: 



6 x 1.5 m x 150 m = 1,350 m 2 . 

Domestic products of 50- by 50-mm mesh 
and 3- or 3.8-mm wire cost $3.45/m 2 or 
$4.73/m 2 , for a total of $4,650 or 
$6,400, respectively. Clips, wire rope, 
and tools add 15 pet of the above to the 
material cost. However, material cost is 
negligible compared to the labor cost for 
installing the mesh. Hence, a more pli- 
able grade of wire mesh, even though more 
expensive, could derive benefits because 
of easier application. 



Revenue shortfall 
Opportunity loss 
Labor 

Total cost 



$152,000 

9,000 

8,000 

169,000 



This total cost does not include equip- 
ment-related interest and depreciation. 
If mechanization could shave off one- 
third of the mesh-application time, the 
monetary gain coupled with the hazard re- 
duction would be sizable. All equipment 
needed for mechanical mesh laying can be 
easily fabricated in a mine shop. 



CHEMICAL ROCK STABILIZATION 



Under the effect of mining, coal and 
adjacent rock tend to break up in blocks 
along planes of weakness such as cleats, 
joints, crevices, cleavages, faults, and 
clay veins. Rock stabilization by chemi- 
cal means can prevent such disintegration 
and keep the rock body intact. Polyure- 
thane has been applied in West German 



mines since the early sixties to accom- 
plish stabilization of fragmented strata 
06). 

Polyurethane has the following prop- 
erties that make it suitable for rock 
stabilization: 



11 



• Low initial viscosity, allowing pen- 
etration into the smallest crevices; 

• High expansion through foaming (al- 
lows it to fill all openings); 

• Controlled rate of increasing vis- 
cosity after mixture (prevents the 
end product from flowing out of the 
stabilization space); 

• Plasticity (so it follows rock move- 
ments) ; and 

• Adherence to rock, coal, wood, ce- 
ment, and steel. 

Polyurethane is a two-component resin 
applied either in cartridge form or by 
injection. Cartridges contain the two 
reactive components separated by plas- 
tic tubing. They are 0.3 m long and 
are inserted into 50-mm-diam boreholes 
at a ratio of approximately 1.5 car- 
tridges per meter of borehole. For in- 
stance, a borehole 3 m long will require 
5 cartridges. 

After the cartridges are pushed to the 
end of the hole, a hardwood dowell is in- 
serted into the hole and rotated by 
the drill for 20 to 30 s to destroy the 
cartridges , mix the two components , and 
thus, to initiate the reaction. The 
borehole is then sealed. Cartridges are 
mostly used to stabilize coal on longwall 
faces in thick seams , and thus to coun- 
teract the tendency of such faces to hade 
forward, which can leave a wide span of 



unsupported roof between canopy tips and 
the face (fig. 10). 

Polyurethane injection is a more gen- 
eral application, suited to the stabili- 
zation of fragmented roof rock (fig. 
11). The two components of polyurethane, 
polyisocyanate and polyol, react after 
mixture at a ratio of 1:1 by volume. The 
polyisocyanate is a dark-colored fluid. 
Polyol is a clear fluid and is available 
in varieties according to starting, reac- 
tion, and hardening times, as shown in 
table 2. 

Systems with accelerated reaction are 
used in rock strata where the process of 
deterioration has gone so far that the 
standard-type resin would flow out of the 
rock before it hardens. Examples of such 
critical areas might be face ends in 
retreat longwall mining where abutment 
stresses preceding the longwall face com- 
bine with those from the gob area of an 
adjacent mined-out panel. 

Fast and standard systems are often 
used in conjunction. Large cleavages are 
closed quickly by the fast-reacting res- 
in, and the final stabilization takes 
place after the standard resin is added. 
Water-resistant systems develop foam 
quickly to seal the strata against the 
influx of water. 

The chemicals are supplied and trans- 
ported in closed metal containers, either 
19-L cans or 209-L drums that are color 
coded for identification. The resins are 



TABLE 2. - Reactive characteristics of polyurethane systems 
(Mixture ratio, 1:1 by volume) 



Characteristic 


Type of polyol 




Standard 


Fast 


Water resistant 




180- 360 

1,200-1,800 

90- 120 

5:3 


65 

95 

15 

3:2 


45-65 

60-70 

15 




8:1 



'Beginning of mixture to start of foaming. 
2 Beginning of mixture to end of foaming. 
3 Foam-to-f luid ratio, by volume. 



Source: Meyer (16, p. 832). 



12 




|§y ,U M /Cartridge 
borehole 






Coal 



2.5 m 



2.5 m 



shear shear 

FIGURE 10. - Coal face stabilization using resin cartridges. 




mRmmmmmmm 



FIGURE 11. - Roof stabilization by polyurethane injection. 



13 



pumped separately from the drums by com- 
pressed air or hydraulically driven units 
and pushed through a labyrinth-type mixer 
immediately before they enter the bore- 
hole. The mixture is injected into the 
fragmented rock mass under a pressure 
approaching 14,500 kPa. Small pump units 
have been designed for direct use on the 
longwall face, and larger units are in- 
stalled in the gateroads to pump the res- 
ins through hose lines to the scene of 
application. Effective telephone commu- 
nication between the pump attendant in 
the gateroad and the injection operator 
on the face is essential. 

Packers were developed to seal the 
polyurethane in the borehole. A packer 
consists of metal shells and a rubber 
hose section. The metal shells , which 
are spread against the borehole walls 
by a setting tool, hold the packer in 
the hole. The rubber hose section pro- 
vides the sealing function. Packers are 
equipped with check valves to prevent 
resin from flowing out of the packer. 

The volume of the infused resin ex- 
pands 3 to 5 times as a result of foam- 
ing. The roof must be supported ade- 
quately to withstand downward movement of 
the newly consolidated rock mass sub- 
jected to injection and expansion pres- 
sures. The injection pumps stall at 
pressures exceeding 14,500 kPa, and this 
indicates that the borehole will not ac- 
cept any more resin, because the material 
pressed into the crevices has set. 

The array of injection holes is pat- 
terned according to site-specific consid- 
erations. As a rule, the length of bore- 
holes should exceed the daily longwall 
advance by 0.5 to 1.0 m. 

Since its introduction in the seven- 
ties, polyurethane has found many appli- 
cations in U.S. deep mines for the sta- 
bilization of fragmented rock on faces, 
at face ends, during shield recovery, and 
in gateroads, often in faulted zones 
( 17-18) . The U.S. consumption of resins 
amounted to 544,000 kg in 1982, and manu- 
facturers estimate the industry will use 
1,360,000 kg in 1983. The estimate for 



European annual use of resins stands at 
18 million kilograms (19). 

The Mine Safety and Health Administra- 
tion (MSHA) has formulated guidelines for 
the safe use of polyurethane in strata 
consolidation applied in cartridges or by 
injection (20). Spraying of polyurethane 
is prohibited. These guidelines encom- 
pass properties of chemicals, such as 
flashpoint, toxicity, and curing tempera- 
ture; quality assurance; equipment stan- 
dards; and operational procedures. Pumps 
must be provided with pressure relief 
(17,500 kPa). The burst pressure of fit- 
tings and hoses must exceed 70,000 kPa, 
and a packer of adequate holding power 
(52,500 kPa) must be used in the bore- 
hole. Recommended operational procedures 
concern underground storage of allowable 
quantities of the chemicals in closed 
metal containers, potable water for 
flushing, first aid supplies, protective 
clothing and goggles , telephone communi- 
cation, responsible and knowledgeable 
supervision, and fire protection (be- 
cause each component of polyurethane is 
combustible) . 

On the basis of the guidelines, MSHA 
district managers issue permits for poly- 
urethane application for individual mine 
sites. State permits are also required. 
Pennsylvania does not allow the presence 
of any persons downwind from the scene of 
polyurethane injection. 

Rock consolidation by polyurethane is 
expensive; therefore, to justify its ap- 
plication, strata monitoring by observa- 
tion through a borescope and differential 
roof sag measurement should precede the 
infusion. A kilogram of polyurethane 
costs approximately $2.95. The weights 
of the components, polyisocyanate and 
polyol, are 1.20 and 1.0 kg/L, respec- 
tively. The rate of consumption is site 
specific, depending on the fissuration of 
the strata. 

Figure 12 shows the roof stabilization 
procedure that was initiated in a recov- 
ery area next to a longwall main gate 
where a wedge of coal had to be left in 
place due to hazardous roof conditions. 



14 



Random bridge 
boards and 
single props 



Gob 



LEGEND 

o Injection borehole 
7a, 7b Test boreholes 

5 



Gob 



Not to scale 



fWWWWAAflflflS 



Mm 



, Shields 

Ml 



30 

flfl 




Gob 



FIGURE 12. • Area stabilized by polyurethane injection during shield recovery. 



The strata immediately on top of the 
shields was broken into small pieces. 
After 17 holes (including 2 test holes) 
were drilled and injected with polyure- 
thane binder, the rock above the shields 
appeared to be laced with cracks filled 
with the binder, which consolidated the 
rock mass into a solid unit. The removal 
of the shields from under the stabilized 
zone did not pose any problems. 

It took 6,100 kg of resin to consoli- 
date the strata. The work, including 
mobilization, was performed in 10 days 
by a team consisting of an engineer and 
two craftsmen. The approximate cost was 
$30,000, broken down as follows: 



Resin; 6,100 kg x $2.95/kg = 
Labor 

Subsistence and airfare 
Total 



$18,000 

10,000 

2,000 

30,000 



Polyurethane infusion during the termi- 
nal phase of longwall panel extraction 
is an alternative to blanketing fri- 
able roof strata above the shields with 
wire mesh, which is an arduous and labor- 
intensive task (21) . Holes on 3-m cen- 
ters are drilled in 2 rows, one row 6 m 
from the predetermined final position of 
the shield line and the other row into 
the roof of the recovery space. Each 
hole takes 140 to 320 kg of resin. Roof 
over recovery chutes that intersect the 
recovery room generated by the terminal 
position of the face can also be rein- 
forced by polyurethane infusion. Hauling 
pans and roof supports through the chutes 
will accelerate the recovery procedure 
(22). 

Such a salvage effort may consume 
25,000 to 40,000 kg of polyurethane. 
However, if the duration of the recovery 



15 



phase and labor costs can be significant- 
ly shaved by polyurethane infusion, this 
method may be more cost effective than 



the installation of wire mesh. The 
choice of a strata stabilization method 
is a site-specific consideration. 



CONVENTIONAL CAVITY FILLING 



Wherever roof deterioration has pro- 
gressed to the degree that cavities ap- 
pear above and in front of canopies , the 
practiced method of reestablishing roof 
contact is to build cribs on top of 
the canopies (fig. 10). During support 
advance, the cribbing is destroyed and 
must be rebuilt. The common method of 
supporting cavities in front of the 
canopies is to establish a platform by 
drilling holes into the face at roof lev- 
el. Small-section H-beams , steel pipes, 
and rods are inserted into these holes. 
Their rear ends rest on the canopies 
(fig. 10). Above this platform timber 
chocks are built to the roof of the cav- 
ity. The practice of building and re- 
building cribbing is time consuming, 
hence costly, and involves exposure of 
workers to unsupported roof. 

The NCB has reported that fatal and se- 
rious reportable accidents associated 
with cavity filling are on the increase 
in British mines , and it is a salient 
fact that people with a great deal of 



practical experience, notably a number of 
supervisors, have been victims of such 
accidents (23) ♦ 

The number of miners injured in the 
United States while filling voids is un- 
known. The total number of longwall ac- 
cidents reported during 1975-81 was 
4,254. Of these, 675 were ground control 
related and 3 were fatal. Two of the 
fatal accidents were the result of roof 
cavities; the third was classified as 
a machinery accident, but was the di- 
rect consequence of a roof cavity. A 
jointed canopy jackknifed into this cav- 
ity. When the tip of the canopy dropped, 
it struck the portal bridge of a pass- 
ing plow. The impact shifted the chock 
aside and caused fatal injuries to a 
miner who happened to be in the travel- 
ling space. It is believed that many 
accidents in which a roof cavity is a 
contributing factor may not be so identi- 
fied because of the method of reporting 
accidents in effect. 



STRAIN-ABSORBENT MODULES 



The NCB has experimented with strain- 
absorbent modules (SAM's), which are 
wire-basket-type substitutes for wood 
cribbing used in the support of roof 
cavities. Straw bales or brush wood was 
used for filling roof cavities in 
the early days of coal mining, and the 
strain-absorbent feature of these materi- 
als inspired the development of SAM's. 
SAM's are rectangular, cubical, and tri- 
angular shapes , structured from welded 
wire-mesh panels (fig. 13). The panels 
are assembled and reinforced by helical 
steel binders. The assembled structure 
can be further strengthened by inserting 
inner panels. 



Though filling cavities with SAM's does 
not eliminate exposure to unsupported 
roof , British miners claim that the use 
of SAM's reduces the exposure time con- 
siderably. SAM's can be made up on the 
surface and folded flat for transporta- 
tion through the mine to the face. Add- 
ing a few helical binders is all that is 
needed at the place of application. 

Ease of material handling and quick es- 
tablishment of roof contact are the major 
advantages of SAM's versus hardwood crib- 
bing. Hardwood cribbing of 600- by 100- 
by 100-mm blocks to fill a void of 600 
by 600 by 600 mm weighs 65 kg, while a 



16 








3 






t — 1~ -H 









































































■ ■ 




>■■■ ■■ ■: ■- 







FIGURE 13. - Strain absorbent modules (SAM's) 



respective SAM structure composed of six 
panels, each made of a 50- by 50-mm mesh 
of 6-SWG (5-mm) wire, weighs only 14 kg 
(8). The module size or sizes can be 
tailored to suit the space to be filled. 
In many instances, a single module can 
easily fill a cavity. 

Though SAM's cannot sustain the same 
load as hardwood cribs, they are surpris- 
ingly strong. The NCB had load tests 
conducted to evaluate the strength of 
modules. In one of these tests, a 1,400- 
by 600- by 600-mm SAM composed of six 50- 
by 5-mm wire-mesh panels was reinforced 
in the middle of its interior space with 
a seventh panel of the same mesh. The 
SAM was weighted with incremental loads 
of 2.5 kN each (fig. 14). When a final 
load of 30 kN (12 weights) was attained, 
the module did not collapse, but suffered 
a deflection of only 41 mm. However, the 



test had to be discontinued because the 
stack of weights atop the module became 
unstable and threatened to tumble down 
(23) . A SAM of the same dimensions and 
mesh, but without the inner support of a 
seventh mesh panel, could only sustain 
17.5 kN before it collapsed. 



As a result of 
concluded that — 



these tests, the NCB 



• SAM's are capable of withstanding 
heavy loads and controlling strata 
in the top and sides of cavities. 

• SAM's should be strengthened by 
clipping inner and end mesh panels 
to them with helical steel binders 
in sufficient numbers. 

• SAM's should be shaped to conform to 
the cavities. 



17 



• The resistance to load increases if 
the SAM is filled with mine waste 
rock or foam. 



The powered roof supports beneath 
SAM's should be advanced with judi- 
cious caution. 



INFLATABLES 



An inflatable structure is another de- 
vice that can be used to fill roof cav- 
ities. Research in the United Kingdom 
has determined that a pressure of only 7 
kPa is adequate to prevent most roof 
falls, stabilize a roof cavity, and en- 
able the roof supports to advance (8). 
Although 7 kPa may seem low for cav- 
ity stabilization, a study titled "Air- 
Inflatable Temporary Supports for Coal 
Mine Roof," conducted by Hauser Labora- 
tories under Bureau contract H02 10058, 
revealed that 90 pet of the roof falls in 
room-and-pillar mining could be prevented 
by a support pressure of 14 kPa, and a 
pressure of 7 kPa would reduce roof falls 
by 64 pet. 

Similar information resulted from a Bu- 
reau analysis of U.S. coal mining fatal- 
ity statistics during 1966-68. Median 




FIGURE 14. - SAM tested under 30-kN load. 



fall dimensions were computed to be 4.30 
m long by 3 m wide by 0.3 m thick. Of 
the falls with fatal consequences, 75 pet 
were less than 0.6 m thick, and 90 pet 
were less than 1.2m thick. This indi- 
cates that pressures of 7 to 28 kPa, dis- 
tributed over the area, would be adequate 
to control the immediate roof strata. 

The Essen Research Center for Roof Sup- 
port and Rock Mechanics in West Germany 
determined that the height of 87 pet of 
measured roof cavities is less than 40 
cm. The weight of a 40-cm-thick rock 
stratum can be supported by a load den- 
sity of 9.6 kPa (1.4 psi). Therefore, if 
such a load density can be brought to 
bear against the roof strata, it may be 
kept intact most of the time (9). The 
support mechanism appears to consist of 
the combination of small pressures dis- 
tributed over large areas, the inherent 
shear strength of the rock, and keying in 
of the stratum forming a voussoir arch. 

To generate such small pressures , the 
NCB is experimenting with inflatable bags 
of various sizes. The typical shape of 
these inflatables can best be described 
as pillow shaped (fig. 15). Nylon fabric 
coated with either neoprene or polyvinyl 
chloride (PVC) is used in construc- 
tion, and each inflatable is fitted with 
two safety valves and a nonreturn inlet 
valve. 

An underground test was conducted in 
a British mine to determine if inflat- 
ables could control roof cavities as well 
as hardwood cribbing. A neoprene-coated 
nylon inflatable was used; it had unin- 
flated dimensions of 4.6 by 0.9 m and 
weighed only 9 kg. It would take 900 kg 
of hardwood crib blocks (60 by 10 by 10 
cm each) to fill a void the same size as 
the inflatable. Inflation took 7 min, 
using a low-pressure stowing machine. 
(In other tests, hand or foot pumps were 
used. ) 



Nonreturn 
inlet valve 



Safety valves 




•ii:"-- 



FIGURE 15.- Inflatables. 



Inflatables may be vulnerable to tears 
and punctures tbat can make them unre- 
liable. To compensate for loss of air 
pressure, an air regulator has been 
developed that admits enough air from 
the compressed-air source to maintain 
the pressure. During one test, an in- 
flatable was punctured by a wire. A 
miner using a foot pump inflated the 
damaged device in less than 7 min to the 
desired pressure. 

The British inflatables are designed to 
be deflated, retrieved, and reused, to 
reduce the cost with each use. How- 
ever, the recovery of the inflatables is 
fraught with dangers and almost impossi- 
ble on a shield-supported longwall. A 
cheaper substitute for neoprene coating 
may be Mylar 4 polyester film. Both neo- 
prene- and Mylar-coated nylon have been 
tested by the Bureau of Mines for possi- 
ble use in emergency inflatable stoppings 
in metal mines (24) . Both materials were 
used to construct 3.7-m-diam spheres. 
The neoprene sphere weighed 50 kg and 

^Reference to specific products does 
not imply endorsement by the Bureau of 
Mines. 



cost $3,500, and the Mylar sphere weighed 
16 kg and cost $1,300. 

Dunnage bags are a substitute that is 
considerably cheaper than neoprene- or 
Mylar-coated nylon. They are constructed 
from several plies of kraft paper that 
cover a polyethylene bladder. At 56 kPa, 
a dunnage bag is fully inflated and re- 
sembles the nylon inflatable (fig. 15). 
The transportation industry has used dun- 
nage bags for years to stabilize loads 
up to 30,000 kg (25). Figure 16 shows a 
dunnage bag supporting a 1,270-kg vehi- 
cle. However, despite their high load 
capacity, dunnage bags have a low resist- 
ance to rips and tears. Also, the Bureau 
has reported that, when inflated, dun- 
nage bags become very rigid and inelastic 
(24). 

It is believed that at lower pressures 
dunnage bags may be flexible enough to 
conform to the interior of a void. Dun- 
nage bags can expand only to 300 mm; if a 
larger void is to be filled, more than 
one bag must be used, or a noninflat- 
able fill must be added. The cost of a 
dunnage bag with uninflated dimensions of 
1.8 by 2.7 m is approximately $20. 



19 




FIGURE 16. - Dunnage bag inflated to 12 kPa supporting a 1,270-kg vehicle. 



CHEMICAL FOAM 



West German miners have been using 
synthetic foams for the last 25 yr to 
provide airtight sealing of crib lines 
along longwall gobs and fill cavities on 
faces and along roadways. Chemically, 
these foams are urea-formaldehyde and are 
marketed under the trade name Isoschaum. 
The NCB has imported Isoschaum foam from 
West Germany, and its application on 
British longwall faces has been 
successful. 

Isoschaum is produced from two chemi- 
cals, one a resinous solution and the 
other an acid foam agent. The chemicals 
react under air entrainment and generate 
a foam with a high expansion rate, 25:1. 
The two solutions, in correct dosage, 
are pumped into a mixing gun, where com- 
pressed air under 560 kPa pressure is 
added. The pumps can be powered by the 
oil-in-water emulsion of the face hy- 
draulic system, and compressed air is 



supplied either from a main compressed- 
air network or a portable compressor 
(fig. 17). Miners can apply Isoschaum 
remotely through a lance, without expo- 
sure to unsupported ground. Due to its 
high expansion factor, the foam conforms 
to the shape of the void and quickly pro- 
vides strong support of the stratum. 

The procedure for treating cavities on 
a longwall face with Isoschaum involves 
two steps. First, a floor is built above 
the roof supports; then, the remaining 
empty space is filled with foam. The 
floor may consist of joints of ventila- 
tion tubing with the ends tied or pieces 
of brattice cloth sewn together. The 
floor material is placed into the voids 
above the canopies or their forepol- 
ing extensions and then injected with 
foam that hardens quickly (fig. 18). 
Any remaining empty space is filled with 
more foam (fig. 19). The foam exerts a 



20 




FIGURE 17. • Isoschaum foam application. 




FIGURE 18. - Foam-filled vent tubes on base of forepoles and split bars. 



21 




FIGURE 19. - Cavity filling with foam. 

positive pressure on the surface of the 
cavity, stabilizes it, and prevents meth- 
ane accumulation. 

Filling cavities with Isoschaum may 
also mitigate the impact of additional 



roof falls. In one instance in a British 
mine, a cavity 7 m long and 3 m high had 
been filled with Isoschaum. After mining 
resumed, a second fall occurred in the 
same location. Large blocks of sandstone 
dropped out from the roof of the cavity, 
but were suspended in the foam without 
doing any harm (8_) . 

In table 3, the relative properties and 
costs of various materials that could be 
used to fill a 10 m 3 cavity are compared. 
The table shows that synthetic (organic) 
foam costs 23 pet less than cribbing and 
20 pet less than inf latables , but 13 pet 
more than SAM's. 

Ease of application and a relatively 
low cost made Isoschaum the preferred 
material for filling cavities on British 
longwall faces (at one time), and the 
NCB was looking for domestic sources to 
supply the foam. However, for health 
reasons, the NCB reversed its position; 
it no longer favors Isoschaum for this 
application and has discontinued Isos- 
chaum use in British coal mines. 



TABLE 3. - Comparison of various materials for filling a 10-m 3 cavity 





Expan- 


Method of 


Full 
support 


Requirements for 10-m 3 


cavity 






Volume 






Material 


sion 


transport 


to sides 


Weight, 3 


(to 


Relative 


Relative 




factor ' 


along face 2 


of 


kg 


trans- 


exposure 


cost of 








cavity 




port), 3 
m 3 


time 4 


materials 5 


Wood 


3:1 


Conveyor and 


No 


3,130 


3.33 


100 


100 


cribbing. 




manual transport. 














12.5:1 




No 


420 


.79 


10 


68 


Inf latables. 


47:1 


...do 


No 


35 


.22 


Nil 


96 


Organic foam 


25:1 




Yes 


450 


.65 


Nil 


77 


Inorganic 


8:1 




Yes 


1,480 


1.95 


Nil 


135 


foam. 
















Packbind 


2.5:1 




Yes 


6,000 


4.54 


Nil 


188 


grout. 6 

















1 Foam-to-fluid ratio, by volume, 

2 Important accident risk factor in handling material on coal face. 
3 Important risk factor in transportation on surface and underground. 
4 After completing base of girders, pipes, etc.; basis: exposure time for wood 
cribbing = 100. 

5 Basis: cost for wood cribbing = 100. 

6 Many of these data apply to placing pump packing and other grout forms . 



Source: Lewis (23, p. 12). 



22 



Before the NCB reversed its position 
on Isoschaum, air samples were taken in 
British mines to evaluate the health as- 
pect of Isoschaum application. During 
the hardening of urea-formaldehyde foam, 
quantities of formaldehyde are released. 
In some instances, while advancing roof 
supports under cavities filled with Isos- 
chaum, formaldehyde concentrations of 1 
to 2 ppm were detected. Even these low 
concentrations can cause watering and 
burning of the eyes and general upper 
respiratory irritation. 

In consideration of studies performed 
by the Chemical Industry Institute of 
Toxicology (CUT), the U.S. National In- 
stitute of Occupational Safety and Health 
(NIOSH) recommends that formaldehyde be 
handled in the work place as a potential 
occupational carcinogen (26) . The Con- 
sumer Product Safety Commission (CPSC) 
banned the use of urea-formaldehyde as a 
home insulating material because it re- 
ceived numerous complaints concerning ad- 
verse health effects (27) . However, this 
ban was overruled by a Federal court in 
July 1983. On the basis of the NIOSH 
recommendations and the CUT studies, 



MSHA has not permitted the use of urea- 
formaldehyde underground. 

Health hazards associated with organic 
synthetic foam such as Isoschaum could be 
avoided if foam could be generated from 
inorganic chemicals that do not give off 
any harmful vapors. The NCB initiated 
experimentation in this area. Dilute 
aqueous solutions of sodium silicate and 
a hardener composed of magnesium and zinc 
salts were fed with compressed air into a 
gun to generate foam. Only 1 pet of an 
organic surfactant was admixed. The lab- 
oratory results were encouraging. Howev- 
er, the expansion factor was only 8:1, 
much less than 25:1 for the organic syn- 
thetic foam; and the cost is much higher 
because this process is still in the 
experimental stage. Underground trials 
have not been reported to date (8). 

However, there has been some recent 
progress. Visitors to the NCB in 1983 
reported that the Mining Research and De- 
velopment Establishment at Bretby has 
succeeded in developing an aerated cement 
system called Aqualight, a pumped fill 
that expands at a rate of 15:1 (28). 



MONOLITHIC FILLS 



Though roof support by shields offers 
distinct advantages, cavities may occur 
both above and in front of the canopies 
under adverse ground conditions such as — 

• Friable immediate roof strata. 

• Laminated shale that tends to 
exfoliate. 

• Distinct slip and separation planes. 



until firm ground is reached. Ground 
stabilization by inserting resin-anchored 
wood dowels or by polyurethane injection 
has been effective where roof cavities 
have steep flanks and are of little ex- 
tension. However, where the coal face 
is hading forward, leaving a wide unsup- 
ported span between the canopy tips and 
face, weak roof strata may collapse; and 
very high and extended cavities may form 
and become uncontrollable (fig. 20). 



• Wedges of shale thinning out under 
massive strong strata. 

• Erosional channels, often called 
rolls. 

• Main joints paralleling the longwall 
face. 

Shields can overcome shallow cavities 
of little expanse by simply advancing 



To remedy such extremely adverse condi- 
tions, West German and British miners 
have successfully applied strata replace- 
ment techniques by applying monolithic 
packs made of various materials similar 
in composition to those used for road- 
side packs (29) ♦ The major monolithic 
approaches to date are (1) the anhydrite 
system and (2) liquid systems such as 
"pump packing" or "aqua packing." 



23 



Vr % '" * / .. / „ Massive ' , v u « ; ' 

*'•'."" " . %; ^ ^ -.• *:' ./"■;, sandstone. '"^'^f f i^.^^'>s 



• ' v. •■'.*. • 



•'.- J' .\ * - * • 




FIGURE 20. - Cavity over coal face hading forward. 



Anhydrite-type packs are prepared by 
blowing a mixture consisting of natural 
anhydrite, water, and a hardener through 
100- or 125-mm airpipe into a frame built 
with a few timbers and a light netting or 
brattice cloth stretched in between. The 
mixture comes through the pipe slightly 
damp and has an angle of repose 30° off 
the vertical. It sets up rapidly and may 
attain a compressive strength of 7,000 
kPa after 10 h and 14,000 to 21,000 kPa 
after a few days. 

The dry anhydrite is transported into 
the mine in bulk, stored underground in a 
bunker, and fed into a blowing machine 
capable of sending a continuous stream of 
anhydrite through the pipeline to the 
scene of application at a rate of approx- 
imately 6 t/h. The hardener is composed 
of potassium and ferrous sulphates and is 
apportioned to the anhydrite in a quan- 
tity equal to 1 pet of the mixture. The 
anhydrite-to-water ratio is 10:1. 



The pump-packing system, developed in 
1973 in the United Kingdom, utilizes a 
mixture of coal fines, bentonite (flow- 
mat), water, and a type of cement called 
Packbind to form a monolithic pack. For 
remedial application only, the Packbind 
component is used with water. Packbind 
is a specially formulated cement that 
sets up rapidly even though substantial 
amounts of coal may be mixed with it. 
The mixture is a slurry that is pumped to 
the scene of application into strongly 
built shuttering. 

Aqua packing (fig. 21) is a variation 
of pump packing that uses essentially 
the same pumping equipment but dispenses 
with the need for coarse aggregates by 
utilizing increased quantities of water 
and special cements. The resultant mix 
crystallizes and sets to produce a pack 
having properties similar to those of 
a standard pumped pack. As an alterna- 
tive to shuttering, the slurry can be 



24 



Feed line for bentonite-sodium 
carbonate blend 



Packhole 




/ 
Power pack 



/ 

Mixer for 
bentonite-sodium 
carbonate blend 



Feed line for 

Portland cement 

blend 



Power pack 



/ 
Mixer for 
Portland cement 
blend 



FIGURE 21. - Aqua-packing system. 




FIGURE 22. - Strata replacement in a high 
and wide cavity. 

pumped into bags tailored to fit the cav- 
ity. British mining engineers believe 
that the maximum compressive strength of 
pumped pack material, about 8,400 kPa, is 
adequate. 

West German miners successfully applied 
remedial techniques as described in 



reports from the early seventies. One of 
the first longwalls with shields oper- 
ated in the Dickebank Coalbed, extracting 
2.21 m of the seam under a roof of rash 
with several slip planes (9) . Rather 
than move the bulky face equipment from 
panel to panel, the entire longwall was 
swung 180° to form a new face line. Dur- 
ing this rotation the weak roof strata 
dropped out several times. 

The following measures were taken to 
stabilize the ground before resuming 
mining under a 10-m-high cavity (30) 
(fig. 22): 

• Rails were placed into holes drilled 
into the face on 1.5-m centers. The 
rear ends of the rails rested on the 
shield canopies. 

• Bridge boards, laid above the rails, 
provided a tight shuttering. 

• Natural anhydrite was blown into 
the cavity on top of the bridge 
boards to form a cover 1.5 to 2.10 m 
thick. 

In addition, the face was secured with 
resin-anchored wood dowels and polyure- 
thane injections. Without any further 
delay, the cavity was undermined, and, 
after 10 shearer passes, normal operating 
conditions were reestablished. 



25 




Sample brought out, 
possibly silly mudstone 
with ironstone bands 



FIGURE 23. - High cavity on shield face. 



Figure 23 shows a very high cavity over 
a heavy-duty shield support on a British 
face 210 m long and extracting 2.74 m 
of the Barnsley/Dunsil Coalbed (31). A 
lens of mudstone was overlain by a strong 
sandstone. The soft mudstone dropped out 
from under the massive strata, while the 
coal face haded forward. The cavity grew 
progressively worse until it attained a 
height of 18 m and an extension of 15 m 
along the face. To stabilize the fall 
area, management decided to fill the cav- 
ity with Packbind, which is quick, set- 
ting, attains early strength, and can be 
cut by a shearer. 

The stabilization program was conducted 
in the following phases (fig. 24): 

1. Shuttering, consisting of plywood 
panels, was built in sections of 1.5 by 
0.75 m. Ten of these sections formed a 
base, extending over the 15-m length 
of the cavity. Four rows of the 0.75-m 
high sections raised the shuttering to a 
height of 3 m. The shuttering was braced 
with proper spragging. 



Shuttering 




15m 



1.5 m 



PLAN VIEW 



Approx l-m girders 
plated together A 

parallel to face 



Shield tongues extended to 
support girders where possible 




4.5 m 



150- by 75-cm 
sectional shuttering 

SIDE ELEVATION 



FIGURE 24. - Strata replacement scheme 
used to fill cavity shown in figure 23. 



2. Packbind was pumped into the space 
between the coal face and shuttering. 
The Packbind was mixed in a grout mixer 
station located in the return gate (18 m 
from the face line) and pumped through a 
50-mm hose into the shuttered area (110 m 
down the face) . The Packbind-to-water 
ratio was 1:2 by weight. Packbind was 
supplied in 25.4-kg bags. The system was 
cleared with fresh water after every 100 
bags pumped. The armored face conveyor 
was kept running during this procedure to 
prevent clogging of the bottom race with 
Packbind, which leaked from the shutter. 

3. Rails 3 m long were placed over 
the Packbind fill, one over each shield 
canopy. 

4. Another row of shuttering, 1.5 m 
high was erected above the rail level. 

5. Packbind was pumped in again, to 
the top of the shutter, to complete a 
fill 4.5 m high and 15 m long, which re- 
established the face line. 



26 



The entire process, from building the 
shuttering to completing the fill, took 
48 h and required 97 t of Packbind. 

Mining was resumed cautiously. The 
shearer cut through the fill while the 
shields were advanced under brushing con- 
tact with the artificial roof. As the 
coal was reached, normal operation was 
reestablished. 

The major differences between the mono- 
lithic systems can be summarized as fol- 
lows: Anhydrite sets up rapidly, is 
stronger than pumped-pack and aqua-pack 
systems, and is installed with relative 
simplicity, needing only light shutter- 
ing. However, pneumatic placement of 
anhydrite requires a source of compressed 
air either from a mainline net or from a 
compressor station underground. Another 
serious problem that must be addressed is 
dust propagation at the blowing machine 
and during pack installation. However, 



West Germany's Silicosis Research Insti- 
tute has certified that natural anhydrite 
from to 6 mm (the type and size used in 
mines) contains less than 1 pet silica 
and is not considered to be harmful. 

Anhydrite is not widely mined in North 
America. There is a mining operation in 
Nova Scotia, and there are large deposits 
of natural anhydrite in Utah and other 
places. 

Because of differences in the systems 
and differences in the strata at differ- 
ent mines, determination of the best sys- 
tem for strata replacement will be a site 
specific choice. 

An alternative to monolithic fills is 
lightweight aggregate concrete blocks 
such as cinder blocks, stacked up as a 
dry wall and coated with concrete, that 
can be cut by a shearer and removed from 
the face (8). 



PLANNED BUREAU RESEARCH 



There are several areas in which 
planned Bureau research may reduce expo- 
sure of workers to ground hazards and im- 
prove roof control on longwall faces. 

The application of inflatables merits 
attention as to how to fit them into the 
operational cycle and how to lower mate- 
rial costs. The Bureau plans to investi- 
gate cheap versions of inflatables such 
as dunnage bags and their resistance to 
rips, tears, and water. 

Since the application of Isoschaum foam 
has been challenged on health grounds , 
substitution of a material without toxic 
or carcinogenic side effects will be in- 
vestigated. Foam generated from anor- 
ganic constituents will be a candidate. 



Planned studies into the technology of 
monolithic packs will open a new avenue 
in the art of strata replacement, which 
heretofore has been almost unknown in the 
American mining scene. 

The Bureau's research into remedial and 
strata replacement techniques is directed 
toward the long-range goals of enhancing 
safety, productivity, and resource re- 
covery. As mining penetrates to greater 
depths, growing interaction between over- 
mined and undermined coalbeds is expected 
to make room-and-pillar mining so vulner- 
able to ground hazards that extraction 
by longwall techniques will be the only 
viable alternative. 



SUMMARY AND CONCLUSIONS 



Uncontrollable strata problems that 
lead to stoppages on highly productive 
longwall faces can result in sizable 
losses. A longwall face supported by 
shields and extracting coal with a 
double-ended ranging-drum shearer has a 



potential for a daily average output of 
3,120 t. Given the current realization 
of $36.5/t of utility fuel, the revenue 
shortfall is $114,000 for each day of 
stoppage, and the opportunity loss per 
day is $7,000 with $8 million invested 



27 



and a 20-pct return on investment, in 
addition to $6,000 face labor and inter- 
est and depreciation related to the 
equipment. 

Techniques designed to maintain the 
roof intact are wire meshing and chemical 
rock stabilization. Wire mesh laying has 
been mechanized in Great Britain to mini- 
mize miners' exposure to roof hazards. 
Mechanical mesh laying could find appli- 
cation in U.S. coal mines during the sal- 
vaging phase of roof shield operations 
that pose an unusual problem because part 
of the structure is under caved rock. 
Chemical rock stabilization is indicated 
when, in disturbed roof conditions and 
under the effect of mining, coal and ad- 
jacent rock tend to break up in blocks 
along planes of weakness. 

Wherever roof deterioration has pro- 
gressed to the degree that cavities ap- 
pear above and in front of the canopies, 
remedial action is indicated. Remote 
placement of a fill in support of the 
stratum is an alternative to the usual 
practice of placing cribbing by hand and 
thus reduces workers' exposure to haz- 
ardous roof. Alternatives developed in 
Europe that reduce hazardous exposure are 
SAM's, inf latables , chemical foam, and 
strata replacement techniques that use 
monolithic fills. SAM's are placed 
quickly, but their application requires 
brief exposure of crews to the cavity 
roof. The other systems allow remote 
filling of roof cavities , but need a mea- 
sure of preparation, such as shuttering, 
which involves exposure. 



Table 3 shows how various remedial ma- 
terials used in British mines compare 

REFERENCES 



with wood cribbing for filling a 10-m 3 
void. Ease of transportation for each 
material is related to its expansion fac- 
tor. The higher the factor, the easier 
the material is to transport. SAM's can 
be folded flat. Inflatables increase 
enormously in volume. Synthetic foam, 
through air entrainment during its gen- 
eration, greatly expands and exerts 
positive pressure on the surface of the 
cavity. Monolithic fills of anhydrite or 
cementitious nature provide little expan- 
sion, but air-entrainment techniques or 
admixtures of a foaming agent may improve 
their expansion properties. 

The relative values in the comparative 
cost column of table 3 may be valid for 
the U.S. coal industry. In addition to 
material costs, the opportunity loss that 
results from roof deterioration on the 
face must enter into the evaluation of 
cost effectiveness in each site-specific 
case. 

The cost of application reflects the 
relative simplicity of installation. The 
use of most remedial systems requires 
little training. European miners are 
familiar with monolithic fills that are 
pumped or placed pneumatically. However, 
in the Uited States, this technology 
is in its infancy and is in need of 
development. 

The ultimate goal of cavity treatment 
from a sheltered location is best ful- 
filled by a combination of remedial sys- 
tems. SAM's or inflatables can provide 
rapid "first aid" prior to the applica- 
tion of Isoschaum foam or monolithic 
fill. 



1. Kundel, H. Die Strebtechnik im 
Deutschen Steinkohlenbergbau im Jahre 
1982 (Longwall Technology in the German 
Coal Mining Industry in 1982). Gluckauf , 
v. 119, No. 11, 1983, pp. 512-522. 

2. Gross, M. A. 1977 Census of 
Longwall Installations. Off the Wall. 
Huwood-Irwin Co., Pittsburgh, PA, Aug. 
1978, 6 pp. 



3. Gross, M. A. (Dep. Energy.) Pri- 
vate communication, 1980; available upon 
request from E. A. Curth, BuMines , Pitts- 
burgh, PA. 

4. Gross, M. A. Census of Longwall 
Installations. Coal Age, v. 85, Dec. 
1980, pp. 89, 91, 93, 95, 97, 99, 101. 



28 



5. Sprouls , M. W. Longwall Census 
'82. Coal Min. and Process., v. 19, Dec. 
1982, pp. 43-47, 50-53, 56-69. 



Min. and 
pp. 49-51. 



. Longwall Census '83. Coal 

Process., v. 20, Dec. 1983, 



7. Hurck, G. Sicherheitliche Prob- 
leme der zunehmenden Teufe und ihre Be- 
waltigung im Steinkohlenbergbau (Safety 
Problems Associated With Increasing Min- 
ing Depth and Their Solution in the Coal 
Mining Industry). Gluckauf , v. 117, No. 
16, 1981, pp. 1061-1066. 

8. Mining Research and Development 
Establishment. Cavity Filling on Long- 
wall Faces. Coal Face Inf. Bull. CF/ 
81/1, Apr. 1981, 25 pp. 

9. Curth, E. A. Coal Mining Tech- 
niques in the Federal Republic of 
Germany— 1971. BuMines IC 8645, 1974, 
54 pp. 

10. Lubba, A., and K. H. Voss . Draht- 
verzug in Streben mit Schreitausbau (Wire 
Mesh on Longwalls With Powered Roof Sup- 
port) . Gluckauf, v. 109, No. 20, 1973, 
pp. 990-992. 

11. Bell, R. L. The Mechanical Appli- 
cation of Wire Mesh Over Powered Sup- 
ports. MRDE Internal Rep. 80/23, 1980, 
5 pp. 



12. 



The Mechanical Application 



of Wire Mesh Over Powered Supports at 
Bilsthorpe Colliery. MRDE Internal Rep. 
78/21, 1978, 8 pp. 

13. Brooks, N. E. Longwall Moving 
Procedure at Consolidation Coal's Mounds- 
ville Operation. Min. Congr. J., v. 65, 
June 1979, pp. 49-51. 

14. Curth, E. A. Safety Aspects of 
Longwall Mining in the Illinois Coal Ba- 
sin. BuMines IC 8776, 1978, 37 pp. 

15. Adam, R. , R. Pimentel, and 
W. Schoff. A Handbook for Face-to-Face 
Moves of Longwall Equipment. Final 



report on BuMines contract J0333925 with 
Ketron, Inc., Oct. 1982, 116 pp.; avail- 
able upon request from E. A. Curth, Bu- 
Mines Pittsburgh, PA. 

16. Meyer, F. Reactive Kunstharze im 
Bergbau (Reactive Resins in Mining). 
Gluckauf, v. 117, No. 14, 1981, pp. 831- 
835. 

17. Pothini, B. R. , and H. A. Von 
Schonfeldt. Polyurethane Aids Roof Con- 
trol in Longwall Mining at Island Creek 
Coal Company. Ch. in Longwall-Shortwall 
Mining; State-of-the-Art , ed. by R. V. 
Ramani. SME-AIME, 1981, pp. 155-166. 

18. McCabe, K. W. Roklok Polyurethane 
Binder: A Chemical Injection System for 
the Consolidation of Severe In-mine 
Ground Conditions. Paper in Proceedings 
of the 1st Annual Conference on Ground 
Control in Mining, WV Univ., Morgantown, 
WV, July 27-29, 1981. WV Univ., 1981, 
pp. 106-115. 



19. 



(Product Manager, Mobay 



Chemical Corp.). Private communication, 
1983; available upon request from E. A. 
Curth, BuMines, Pittsburgh, PA. 

20. Nagy, J. Guidelines for the Use 
of Polyurethane Binder for Strata Consol- 
idation in Underground Mines Where the 
Ambient Temperature is Less Than 80° F. 
MSHA (Dep. Labor), Pittsburgh, PA, Dec. 
15, 1978, 6 pp. 

21. Stewart, J. G. , and S. G. Young. 
Roklok Polyurethane Binder Systems: New 
Developments for Controlling Roof Strata 
During Longwall Face Changes and Sealing 
Water Out of Underground Workings. Paper 
in Fourteenth Annual Institute for Mine 
Safety and Health Research, VA Polytech. 
Inst., Blacksburg, VA, Aug. 23-25, 1983, 
VA Polytech. Inst., 1983, pp. 177-183. 

22. Curth, E. A. Longwall Mining of 
Thin Seams . Paper in Proceedings of the 
1st Annual Conference on Ground Control 
in Mining, WV Univ., Morgantown, WV, July 
27-29, 1981. WVUniv., 1981, pp. 239- 
259. 



29 



23. Lewis, S. , and L. R. Stace. 
Strata Substitution and Reinforcement 
Techniques in the United Kingdom. Paper 
in Seventh International Strata Confer- 
ence, Liege, Belgium, Sept. 20-24, 1982. 
INIEX, 1982, 18 pp. 

24. Thimons, E. , and F. Kissel. An 
Evaluation of Emergency Inflatable Stop- 
pings for Use in Metal Mine Fire Rescue 
and Recovery. BuMines RI 8162, 1976, 
13 pp. 

25. Nocek, R. S. Air Bags - What You 
Need To Know About Selection and Applica- 
tion. Pulp and Paper, v. 52, June 1978, 
4 pp. 

26. Blackwell, M. , H. Kang, A. Thomas, 
and P. Infante. Formaldehyde: Evidence 
of Carcinogenicity. J. Am. Ind. Hyg. 
Assoc, v. 42, July 1981, pp. A-34 to 
A-42. 

27. Halmos, E. E. Safety Commission 
Bans Formaldehyde; Challenge Appears Cer- 
tain. Air Cond. , Heat, and Refrig. News, 
v. 155, No. 9, pp. 1, 4. 



28. King, R. (BuMines Pittsburgh Re- 
search Center) Private communication, 
1983; available upon request from E. A. 
Curth, BuMines Pittsburgh, PA. 

29. Hill, F. E., C. Peake, and 
A. Sharkey. An Analysis of the Applica- 
bility and Probable Cost Effectiveness of 
Advancing Longwall Mining Systems in the 
United States. Final report on BuMines 
contract J0123924 with Emory Ayers Asso- 
ciates, Inc., June, 1981, 146 pp.; avail- 
able upon request from E. A. Curth, Bu- 
Mines, Pittsburgh, PA. 

30. Grundmann, J. Neue Erfahrungen 
mit dem Schildausbau (Recent Experience 
With Roof Shields). Gluckauf , v. 109, 
No. 3, 1973, pp. 198-201. 

31. Dutton, B. A Successful Example 
of the Strata Replacement Technique. 
Min. Eng. (London), v. 140, Apr. 1981, 
pp. 733-739. 



ftU.S. CPO: 1985-505-019/20,023 



INT.-BU.OF MINES, PGH., PA. 27933 



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